1. Field of the Invention
This invention relates to the smelting of tin-bearing materials, and in particular to the production of a tin-bearing gas which is separated from molten slag.
2. Description of the Prior Art
Conventional treatment of tin-bearing materials usually involves two stages. The first includes mixing tin ore or concentrate with slag-forming fluxes (such as limestone and sand) and a reducing agent (such as carbon). The mixture also sometimes includes various by-products, such as fume or dust, dross, and hardhead, which is an alloy of tin and iron, typically containing less than 60% iron because of the relatively high melting point of tin-iron alloys of higher iron content.
The mixture is fed to a smelting furnace, usually a reverberatory furnace, and is smelted to produce liquid metallic tin and a molten silicate slag. This stage of the treatment is subject to the physico-chemical equilibrium between metallic tin and metallic iron dissolved in that tin, on the one hand, and the lower oxides of tin and iron (SnO and FeO, respectively) mixed with other oxides, including silica in the slag, on the other hand. This equilibrium demands that if the iron content of the metallic tin is to be controlled at a reasonable level, typically 1% iron by weight or less, then the tin and iron content of the slag shall be approximately equal. Since the ultimate outlet for the iron is in the slag, the iron content of the slag must be equal to that of the raw material smelted, plus that of the hardhead added to the charge, and since the slag must have a tin content approximately equal to the iron content of the slag (hereinafter referred to as "rich slag"), the slag will typically contain between 8% and 30% tin by weight.
After the smelting furnace charge has been heated to the required temperature to smelt the mixture and produce a lower layer of molten tin metal (plus certain metal impurities which may be present) and a top layer of molten rich slag, the contents of the smelting furnace are tapped out through a siphon or similar device to separate the molten tin metal from the rich slag. The stream of slag is either allowed to freeze on a flat surface so it can be subsequently broken up, or the stream of slag is granulated by a powerful jet of water.
This concludes the first stage of the conventional treatment, which is followed by a second-stage treatment to recover the tin in the rich slag, which is either in the form of dry lumps that must be crushed to suitable size, or else is in the form of wet granulated material of the desired size, and must be dried.
The properly conditioned rich slag is mixed with fluxes and carbon and is charged to a furnace where it is smelted to form hardhead and a second slag (hereinafter referred to as "poor slag").
The high melting point of hardhead precludes the use of a siphon or similar device for the separation of hardhead from the poor slag. Accordingly, separation of the molten hardhead and poor slag must be effected by decantation from open vessels. This process is laborious and inefficient, because some poor slag must be left with the hardhead to prevent loss of hardhead when the poor slag is discarded.
The hardhead must be granulated by a water jet, because it cannot economically be reduced to a suitable size with a crusher. The metallurgical requirement at this stage is that the poor slag shall contain a weight of iron equal to that in the concentrates and other new material (not by-products) fed to the first stage of the recovery operation.
The physico-chemical equilibrium requires that the tin content of the poor slag be about one-tenth of the iron content. If the iron content is so high that the tin content is unacceptably high, a third stage of treatment may be required.
The conventional two-stage process just described has five principal disadvantages.
1. The tin content of the poor slag, which is the waste product of the process, must contain a tin content approximately equal to one-tenth of the iron content of the raw materials smelted.
2. The tin which is recovered from the rich slag is in the form of hardhead and, therefore, is accompanied by an approximately equal weight of iron, which must be absorbed in the next bath of rich slag which is produced. The absorption of this cycling load of iron into the rich slag may raise the iron content to a level which is highly corrosive to the refractory lining of the smelting furnace. Moreover, since it is necessary that the tin content of the rich slag be approximately equal to the iron content, the resulting tin content of the rich slag places a heavy burden of metallurgical work on the second stage of the process.
3. To avoid loss of the hardhead which contains the tin recovered from the rich slag, it is necessary to return some poor slag with the hardhead. The return of this waste to the circuit increases the waste material content of the raw material being processed.
4. The combined effects of the three foregoing disadvantages limits the effective application of the conventional two-stage process to the treatment of raw materials having a tin content in excess of 50% and an iron content less than 5%.
5. The rich slag must be tapped out of the smelting furnace, solidified, and then be sized (by breaking up lumps of solidified slag) or be discharged as a stream of molten slag in a powerful jet of water to produce granulated slag, so the slag can be adequately mixed with the necessary fluxes and carbon for return to the smelting furnace. These operations cause not only the loss of the heat content of the liquid slag, but also incur additional handling costs for the solidified slag.
The above disadvantages all stem from the attempt to recover the tin by means of a liquid-liquid separation (liquid tin alloy-liquid slag). The disadvantages may be overcome if the tin is recovered by using a gas-liquid or gas-solid separation process involving one or more of the volatile tin compounds, such as stannous oxide, stannous sulfide, or stannous chloride.
Many patents have issued describing various gas-liquid or gas-solid processes for separating tin from iron-bearing materials.
British Pat. No. 1,332,726 discloses a process in which the tin-bearing material is melted and raised to the very high temperature of at least 1500.degree. C. in an electric furnace to volatilize stannous oxide. The volatilization is aided by blowing air across the surface of the melt, which is kept quiescent to minimize the attack of the hot melt on the lining of the container. This process has a high energy consumption due to the temperatures at which the reaction is carried out. Moreover, the rate of volatilization of stannous oxide is slow because of the quiescent state of the molten slag.
The separation of tin in the form of stannous sulfide is accomplished by the well-known "fuming" process in which a sulfur-bearing substance, usually iron pyrites, is added to a liquid slag bath, which is both heated and stirred by oil or coal burners submerged below the surface of the molten slag.
The fuming process was developed for the recovery of tin from poor slags containing typically 3% tin by weight, but it can be, and has been, employed for the recovery of tin from rich slags.
The kinetics of the reaction between tin in a slag and a metallic sulfide dispersed or dissolved in the slag appear to be of the first order, i.e., the rate at which volatile tin compounds are formed depends only on the tin content of the slag undergoing treatment. Thus, the recovery of tin from poor slags is slow, and the fuming process must be operated as a batch process. Moreover, it may be necessary to install a heated holding furnace for the accumulation of slag awaiting treatment.
When the fuming process is employed for the treatment of rich slags, the evolution of tin-bearing fume in the early part of the batch treatment is intense. Accordingly, very large gas filtration equipment must be installed to avoid losses of tin fume. The scale of this filtration equipment greatly exceeds the capacity required during the latter part of the batch operation when the tin content of the slag under treatment reaches the level of a typical poor slag. Thus, a high capital investment is involved for gas filtration equipment far in excess of the capacity which would be required if the fume evolution were at a constant rate.
British Pat. No. 1,337,270 discloses a method which tries to overcome the disadvantages of the batch treatment by feeding finely-ground slag, together with pyrite fuel and air, into a cyclone furnace operated at about 1350.degree. C. to achieve a steady volatilization of stannous sulfide at a uniform rate. But to operate this process in conjunction with the conventional production of liquid-rich slag, the slag must be tapped from the furnace and frozen, thereby sacrificing its heat content. Extra work is also entailed in grinding the solidified slag to the required particle size of less than 3 mm.
A disadvantage of all processes relying on the generation of stannous sulfide is that the entire sulfur content of the materials used is converted to sulfur dioxide, which accompanies the volatile tin compounds in the furnace off-gases, but is not trapped by the gas filtration equipment which collects the tin oxide.
The escape of the sulfur dioxide is an environmental hazard, and the installation of scrubbing plants for the removal of the sulfur dioxide from the off-gas is expensive. Furthermore, the operation of scrubbing plants is costly, and the sulfur-bearing products from such plants may be highly acidic sludges which require special disposal to avoid pollution of water sources.
Processes relying on the volatilization of stannous chloride have been described in a number of patents. For example, Netherlands Pat. No. 2,062 (1919) discloses the use of sodium chloride to form volatile stannous chloride, calcium chloride is disclosed in U.S. Pat. No. 1,931,944 (1934), and zinc chloride and ferrous chloride are disclosed in U.S. Pat. No. 1,843,060 (1932).
More recent patents have referred to the need to recover and reuse the reagents and have paid particular attention to the use of calcium chloride. For example, in British Pat. No. 1,095,122, a pelletized mixture of tin-bearing material with calcium chloride and coke is roasted in a closed retort to form stannous chloride, which is condensed in the cold end of the retort. The stannous chloride is then heated with lime and carbon to form metallic tin and calcium chloride, which may be used again to treat new tin-bearing material.
In Japanese Pat. No. 21,415 (1969), a mixture of tin-bearing material with calcium chloride is heated in a retort at about 980.degree. C., and the resulting gases are passed through cold water to dissolve the chlorides of tin, iron, and other metals. The solution is partially neutralized to remove impurities. The iron content is reduced to the ferrous state, and tin oxide is precipitated by adjusting the pH of the solution to between 3 and 4.
French Pat. No. 2,010,021, British Pat. No. 1,266,711, and Canadian Pat. No. 907,865 (all by the same inventors), propose that a mixture of tin-bearing material and calcium chloride be pelletized and treated in the presence of coal at about 900.degree. C. to 1050.degree. C. in a rotary kiln heated by shell burners. The patents state that for a satisfactory kiln process, one must not exceed the sintering temperature of the charge. The issuing gases are passed through a scrubber, and the solution is neutralized with chalk or milk of lime in a step-wise manner to precipitate sequentially impurities, such as arsenic, antimony, and lead, then tin oxide, and finally iron and zinc hydroxides. Each of the three fractions is filtered off before the next neutralization step, and the final clean solution of the calcium chloride is evaporated for reuse. The patents state that in the preferred embodiment, the solution fed to the scrubber is maintained in a highly acid condition to avoid premature precipitation of stannous hydroxide or stannous oxychloride.
The foregoing prior art processes relating to gas-solid separation have in common the need for sophisticated methods of charge preparation, such as pelletizing or briquetting, and also for the careful control of charge temperature in the treatment furnace to prevent sintering of the mass, which must be kept as close as practical to the sintering temperature to achieve a reasonable rate of evaporation of stannous chloride from the solid mass. Thus, the temperature must be carefully controlled so it does not vary more than about 25.degree. C. Even when this is done, the rate of production of volatile stannous chloride gases is relatively slow.